Method for the recovery of gold

ABSTRACT

The invention relates to a method for the recovery of gold in connection with the hydrometallurgical production of copper from a residue or intermediate product containing sulphur and iron generated in the leaching of copper raw material. The recovery of both copper and gold takes place in a chloride milieu. The gold contained in the residue or intermediate product is leached using bivalent copper and oxygen in copper (II) chloride-sodium chloride solution in the conditions, where the oxidation-reduction potential is a maximum of 650 mV and the pH at least 1. The iron and sulphur contained in the residue remain for the most part undissolved.

The invention relates to a method for the recovery of gold in connectionwith the hydrometallurgical production of copper from a residue orintermediate product containing sulphur and iron generated in theleaching of a copper raw material. The recovery of both copper and goldtakes place in a chloride milieu. The gold contained in the residue orintermediate product is leached using bivalent copper and oxygen in acopper (II) chloride-sodium chloride solution in an environment, wherethe oxidation-reduction potential is a maximum of 650 mV and the pH atleast 1. The iron and sulphur contained in the residue remain for themost part undissolved.

Some methods are known in the prior art, which are used for leachinggold from sulphur- and iron-containing material in connection with achloride-based copper recovery process.

U.S. Pat. No. 4,551,213 describes a method whereby gold can be leachedfrom sulphur-containing materials, particularly from the residue ofhydrometallurgical processes. A beneficial source material for themethod is residue from the CLEAR process. The CLEAR process is ahydrometallurgical copper recovery process, which takes place in achloride milieu and at raised pressure. Gold-containing residue isslurried in water and the chloride content of the suspension obtained isadjusted so that it contains 12-38 weight percent of chloride. Theoxygen reduction potential is regulated within the range of 650-750 mVand the pH below 0. Copper (II) chloride or ferric chloride is added tothe suspension to oxidise the gold contained in the raw material so thatthe gold dissolves. The publication mentions that theoxidation-reduction potential must not rise above 750 mV, becausesulphur will dissolve above this value. There is no information on theamount of dissolved sulphur or iron in the publication.

EP patent 646185 relates to copper recovery from sulphidic concentratesusing chloride leaching in atmospheric conditions. Gold is leached fromthe leaching residue into an electrolyte that contains at least twohalides, such as sodium chloride and sodium bromide. The purpose is tostore oxidising power for the bromine complex on the copper electrolysisanode, and use it to leach the gold in the residue.

There are some drawbacks in the methods mentioned above. The leachingconditions in the method of U.S. Pat. No. 4,551,213 are very harsh. Thepatent mentions that sulphur still will not dissolve under theconditions of the patent, but this is not universally applicable sincethe dissolving tendencies of elemental sulphur and the iron compoundsmentioned in the patent depend on the generating method of the sulphurand said compounds. Tests we have carried out have shown that whenleaching residues formed in atmospheric conditions are treated under theconditions in the above-mentioned patent, there is considerabledissolution of sulphur and iron. Obviously their dissolving affects theeconomy of the process. The gold leaching method used in EP patent646185 using a bromine complex on the other hand is not advantageousfrom an environmental viewpoint, since harmful bromine emissions may begenerated in the concentrate leaching stages.

Now a new method has been developed for the leaching of gold from aleaching residue or intermediate product containing iron and sulphur,which have been generated in the atmospheric chloride leaching of coppersulphide concentrate. We have found that it is possible to leach goldfrom an iron- and sulphur-containing material into an aqueous solutionof copper (II) chloride-sodium chloride when oxygen-containing gas isfed into the solution. Leaching takes place thus by means of bivalentcopper and oxygen in conditions where the oxidation reduction potentialis below 650 mV and the pH of the solution is in the range of 1-3. Theoperating range according to this method is clearly more beneficial thanthat mentioned in the prior art, because iron will not yet dissolve inthese conditions and sulphur remains for the most part undissolved. Thisavoids the costs that arise from removing iron and sulphur from thesolution. Leaching occurs in atmospheric conditions at a temperature inthe range between room temperature and the boiling point of thesuspension, preferably however between 80° C. and the boiling point ofthe suspension. Recovery of gold from the solution is made using somemethod of the prior art such as electrolysis or with active carbon. Theremaining sediment is disposable waste.

The essential features of the invention will be made apparent in theattached claims.

A residue or intermediate product containing gold is pulped into asodium chloride solution containing copper (II) chloride to form asuspension and the oxidation reduction potential required for goldleaching is obtained using specifically bivalent copper and oxygen. Theoxidation-reduction potential is measured with Pt and Ag/AgCl electrodesand the potential is held at a value below 650 mV, preferably a maximumof 620 mV. When the oxidation-reduction potential is held below a valueof 650 mV, sulphur will not dissolve from the residue. The preferred pHrange is 1.5-2.5. Below a pH value of 1 the iron in the solids willstart to dissolve, and this is undesirable. Air, oxygen-enriched air oroxygen can be used as the oxidising gas. The amount of bivalent copper,Cu²⁺, in the solution is preferably 40-80 g/l and the amount of sodiumchloride in the range of 200-330 g/l.

It is beneficial to link the method as a sub-process of a copperconcentrate chloride leaching process. A method of this type isdescribed in U.S. Pat. No. 6,007,600. In the said method, a coppersulphide-containing raw material such as a concentrate is leachedcounter-currently with a sodium chloride-copper (II) chloride solution,NaCl—CuCl₂, in several stages in order to form a monovalent copper (I)chloride solution, CuCl. A residue remains in leaching, which containsmainly the sulphur and iron of the raw material as well as the goldcontained in the raw material. The method now developed relates to theleaching of gold from the residue formed in the type of processesdescribed above.

The method of the present invention is further described in theflowchart of FIG. 1, where gold recovery is connected to thehydrometallurgical recovery of copper. The flowchart represents oneexample of an embodiment of our invention. The thicker arrows in FIG. 1show the movement of the solids and the thinner arrows show the flow ofthe solution.

A sulphidic raw material of copper such as copper sulphide concentrate 1is fed into the leaching reactor 2 of the first leaching stage, intowhich is also recirculated solution 3 from a later process stage, whichis an aqueous solution of copper (II) chloride-sodium chloride. Thethicker arrows indicate the flow of the solids and the thinner arrowsthe flow of the solution. The copper from the copper concentratedissolves into the process solution, and the solution is routed tothickening 4. After thickening the overflow 5 contains copper chloride,having about 70 g/l copper mainly as monovalent form, and it is routedto the copper recovery process (not shown in detail). The leaching ofthe solids contained in underflow 6 is continued further in reactors 7and 8 of the second leaching stage with solution 9, which is obtainedfrom a later process stage. The Cu²⁺ content of the solution 9 going tothe second leaching stage is adjusted to its optimum with an NaClsolution. Air is introduced to the reactors of this stage in order tointensify leaching. Thickening 10 is at the end of the stage.

The overflow 3 from thickening 10 of the second stage is routed to thefirst stage to leach the concentrate. The leaching of the solids of theunderflow 11 is continued in the third stage in reactors 12, 13 and 14in order to leach the rest of the copper and the gold. Please note thatthe number of reactors in the flow sheet does not limit the number ofreactors in the method of the present invention. In the third leachingstage i.e. the gold leaching stage, the residue is leached with a strongsolution of copper (II) chloride-sodium chloride 15, where the Cu²⁺content is 60-100 g/l and the sodium chloride content 200-330 g/l.Oxygen is fed into the reactors preferably in the form of air. As theleaching stage ends the slurry is routed to thickening 16. The overflow17 from thickening is routed either as it is or filtered to goldrecovery, which in this embodiment takes place in carbon columns 18using active carbon. The gold product 19 is obtained from the columns.The solution removed from the columns is a gold-free solution 9, whichis recirculated to the second stage of the leaching and if requiredsodium chloride solution 20 is fed into it in order to get a suitablecopper (II) chloride content for leaching. The underflow or residue fromthe gold recovery stage thickening, after normal post treatment such asfiltration and washing (not shown in detail) becomes the final waste 21,which contains almost all the sulphur and iron of the concentrate. Theresidue filtrate and rinse waters are returned for instance to theconcentrate leaching process.

The multi-stage leaching of the copper raw material is shown in the flowsheet as counter-current leaching and within the stage the solid matterand solution move basically uniformly from one reactor to another. Inorder to intensify leaching however, the solids could be recirculated byreturning them within the process. Thus the solids may be returnedwithin one of the stages comprising several reactors, from the tail endreactors to the front end reactor of the stage, or recirculation couldeven be implemented within an individual reactor. At the end of everystage or after each reactor the separation of liquid and solids takesplace, generally using a thickener. The solution obtained from theseparation between stages, that is the overflow, is routed to theprevious stage in the direction of the solids flow and the solidresidue, or underflow, is mainly routed to the following leaching stage.Thus part of the underflow of one or each stage can be returned to areactor from either the previous or the same leaching stage, preferablyto the first reactor.

The flow sheet in FIG. 1 presents a gold leaching method in connectionwith leaching of a copper-containing raw material, but the method of thepresent invention is not bound exactly to the copper-containing rawmaterial leaching process in the flow sheet. The key point in our methodis that the leaching of gold-containing material is performed withbivalent copper and oxidizing gas in conditions where the potential ofthe solution is less than 650 mV, preferably between 530-620 mV and thepH is at least value of 1, preferably at least a value of between1.5-2.5. When the oxidizing gas is air, the reactor structures can beformed simply.

The invention is further illustrated in the following example.

EXAMPLE 1

Conditions according to the prior art (U.S. Pat. No. 4,551,213) for therecovery of gold were used in the example. The leaching residue used inthe tests originated from chloride-based leaching of a copper sulphideconcentrate, performed in atmospheric conditions. The moisture of theresidue was 31% by weight and included 3.7% Cu, 28.9% Fe, 32.4% S and5.8 ppm Au measured as dry weight.

220 g of moist leach residue was placed in a mixing reactor with 500 mlof a solution that contained 40 g/l of Cu²⁺ as chloride and about 300g/l of NaCl. The solution temperature was 40° C. and the leaching time12 hours. During the leaching time the oxidation-reduction potential ofthe slurry in the reactor was kept at a standard value of 680 mV usingchlorine gas, when measured with Pt and Ag/AgCl electrodes. The pH ofthe slurry was allowed to change freely during the test from theoriginal value of 2 to its final value of 0.1. At the end of the testthe analyses of the solution and solids were as follows:

Solution Fe g/l S g/l Au mg/l 42.6 9.33 1.28 Solids Fe % S % Au ppm 19.746.4 3.1

The test results show that about half of the iron dissolved, which wouldcause very great removal costs in a production plant. Only about half ofthe gold dissolved.

EXAMPLE 2

This example was carried out according to the method of the invention. Acopper sulphide concentrate (CuFeS₂) was leached with a CuCl₂—NaClsolution and air in a mixing reactor so that a leaching residue wasgenerated with the following contents (measured as dry weight):

Cu % Fe % S % Au ppm 0.7 41.6 28.6 3.9

The original concentrate contained about 6.8 ppm of gold, and thus partof it had already dissolved when the concentrate was leached. After thisslurry was made from the residue and a new solution, which contained 87g/l of leaching residue and the original solution, which contained:

Cu g/l Fe g/l S g/l Au mg/l 71.2 0.08 0.553 0.016

The copper of the original solution was in cupric form. The slurry washeld in a mixing reactor equipped with a 5-liter air feed for 12 hoursat a temperature of 100° C. The leaching process is illustrated by thefollowing measurements and analysis results, shown in Table 1.

TABLE 1 Solution Residue Time Au Fe S Au (h) pH (reactor) Potential (mV)mg/l g/l g/l Ppm 0 2.4 552 0.016 0.008 0.53 3.9 4 2.7 572 0.136 0.0020.86 8 2.6 610 0.240 0.003 1.09 12 2.6 608 0.260 0.005 1.21 0.8Potential: Pt vs. Ag/AgCl

400 ml of the final test filtrate was taken and 10 g of active carbonwith an average grain size of 1.5 mm. was added, and it was then mixedfor 4 hours at a temperature of 25° C. At the end of mixing the solutionwas analysed and shown to contain <0.005 mg/l Au.

The analysis results show that the iron remained in insoluble form andthat the sulphur also only dissolved a little i.e. about 0.7 g/l. Eventhough the original gold content was low, the leaching yield wasnevertheless a good one, at about 80%.

1. A method for the recovery of gold from a leaching residue orintermediate product containing iron and sulphur, which is generated inthe chloride leaching of a copper sulphide raw material at atmosphericpressure, comprising leaching the gold from the residue or intermediateproduct in an aqueous solution consisting essentially of copper (II)chloride, sodium chloride and oxygen-containing gas; keeping theoxidation-reduction potential of the suspension formed at a value below650 mV and the pH at a value of 1-3, whereby the iron and sulphur remainmainly undissolved; recovering the dissolved gold, and; discarding theundissolved residue as waste.
 2. The method according to claim 1,wherein the oxidation-reduction potential is kept in the range of530-620 mV.
 3. The method according to claim 1, wherein the pH of thesuspension is kept at a value of 1.5-2.5.
 4. The method according toclaim 1, wherein the amount of bivalent copper in the suspension is40-100 g/L.
 5. The method according to claim 1, wherein the amount ofsodium chloride in the suspension is 200-330 g/L.
 6. The methodaccording to claim 1, wherein the temperature of the suspension is keptin the range between 80° C. and the boiling point of the suspension. 7.The method according to claim 1, wherein the oxygen-containing gas isair.
 8. The method according to claim 1, wherein the oxygen-containinggas is oxygen-enriched air.
 9. The method according to claim 1, whereinthe oxygen-containing gas is oxygen.
 10. The method according to claim1, wherein the dissolved gold is recovered using active carbon.
 11. Themethod according to claim 1, wherein the dissolved gold is recovered byelectrolysis.